Synthetic rutile production

ABSTRACT

Hot metal is circulated by an R-H unit in a closed loop path through first and second hearths and chambers. Titaniferous material containing iron oxide is introduced into the first hearth and the iron oxide therein is reduced in a heating zone in the first hearth to iron to produce titania slag having a reduced iron content which is removed in chamber before the hot metal passes via an underflow weir into chamber at which coal is added and a proportion of hot metal is removed. The addition of coal is such that more than 2% by weight of carbon dissolves in the hot metal in the second hearth. Coal ash slag is removed at chamber, while the hot metal containing the dissolved carbon is recirculated to the first hearth.

This invention relates to the production of synthetic rutile of aquality suitable for use as a feedstock in the production of TiO₂pigment via the chloride route.

The chloride route to TiO₂ pigment production has become increasinglyimportant in view of environmental concerns about the acid wastesproduced by the sulphate TiO₂ process which can accept lower gradetitaniferous ores.

Commercial synthetic rutile production for use as a feedstock in thechloride route is normally based on ilmenite feedstock containing 54 to60% TiO₂, with a distinct preference for material of at least 58% TiO₂content.

Various proposals have previously been made for processing lower gradetitaniferous materials as alternative sources of TiO₂ pigment. Forexample, a process has been used which involves blending ilmenite withcoke and smelting in an electric arc furnace to separate most of theiron as saleable pig iron and to form a titania-enriched slag.

Enrichment of the TiO₂ content of conventional titanium slags has alsobeen disclosed by dc arc smelting.

EP-A-0266975 (corresponding to U.S. Pat. No. 4,701,217) discloses amethod of smelting metal oxide material comprising forcibly circulatinga molten carrier material in a closed loop path serially through asmelting reduction zone, a slag separation zone and a heating zone;contacting the metal oxide material with the molten carrier material;introducing a carbonaceous reductant to the molten carrier material; atleast partially reducing the metal oxide to metal by the carbonaceousreductant in the smelting reduction zone, the metal oxide andcarbonaceous reductant being utilised in proportions such that thecarbon from the carbonaceous reductant is converted to carbon monoxide;reacting the carbon monoxide with oxygen in the heating zone at thesurface of the molten carrier material so that the heat generated by thereaction is transferred to the molten carrier material which iscirculated to the smelting reduction zone; separating slag from saidmolten carrier material in said slag separation zone before the moltencarrier material is circulated to the heating zone so that the surfaceof the molten carrier material which is circulated to the heating zoneis substantially free of slag; and recovering said metal. Amongst alarge number of possible feedstocks which are described, EP-A-0266975discloses the application of the above process to ilmenite to formtitaniferous slag (a source of synthetic rutile) and pig iron, but thereis no detailed example of this and no proposals as to how to upgrade thequality of the titaniferous slag to the quality required for TiO₂pigment production via the chloride route.

According to the present invention, there is provided a method ofupgrading titaniferous material containing iron oxide, comprisingcontacting the titaniferous material with molten iron containingdissolved carbon so as to reduce at least some of the iron oxide in thematerial to iron and produce a titania slag having a lower iron oxidecontent.

Most preferably the method of upgrading titaniferous material containingiron oxide, comprises the steps of circulating a molten carrier materialcomprising molten iron and dissolved carbon in a closed loop paththrough first and second hearths; introducing the titaniferous materialinto the carrier material in a heating zone in the first hearth so as toreduce iron oxide to iron and produce a titania slag having a lower ironcontent; removing the titania slag from the molten carrier materialbefore the latter is circulated to the second hearth; introducingcarbonaceous reductant into the molten carrier material so as to causecarbon to dissolve in the carrier material in the second hearth; andperforming a further slag removal operation on the circulating moltencarrier material before passing the latter to the first hearth.

In the above described method, the further slag removal operationeffected in the second hearth after introduction of the carbonaceousreductant serves to remove slag resulting from addition of thecarbonaceous reductant so that such slag does not become mixed with thehigh grade titanium dioxide slag which has formed in the first hearth.Thus, it is possible to use a relatively low grade carbonaceousreductant such as ordinary coal rather than having to use an expensivevery low ash coal. Because the coal ash slag is kept separate from thehigh grade titanium dioxide slag, if necessary, flux materials such aslimestone or dolomite can be blended with the added coal to permitutilisation of coal with associated ash having a relatively high ashfusion temperature without introducing highly undesirable calcium ormagnesium oxides into the titania slag product. This feature isparticularly significant if the circulating molten carrier material isto be maintained at temperatures considerably lower than that of thetitania slag product as will be mentioned hereinafter, and in such casesit is sometimes necessary to ensure adequate fluidity of the coal ashslag by addition of appropriate fluxing materials in the first hearth.

It is highly preferred to control the amount of carbon in the ironrelatively close to carbon saturation, typically so that the carbon isat least 2% and most preferably in the range 3 to 5% by weight. Bycontrolling the carbon at a relatively high level, the kinetics of thesmelting reduction reactions are enhanced, which because of their largeendothermicity tend to keep the solid charge material below its meltingpoint for as long as possible. It has been observed that reductionreactions of solid pellets of ilmenite proceed most rapidly whilst thecharge pellets are solid and floating on the surface of the moltencarrier material. For the same reason, it is preferable not to preheatto an elevated temperature the charge material which ideally ispelletised and sintered to form hard strong and dense pellets, typicallyabout 1.5 cm in diameter. Preferably, the pellets are stored cold inpreparation to being charged carefully to the first hearth so that theyare dispersed uniformly across the width of the molten carrier materialand then float away from the charge area so that initially some 55 toabout 70 percent of the free surface of molten carrier material iscovered with these single charge pellets. Some initial sticking togetherof freely floating pellets is of no consequence because as reactionproceeds they normally tend to separate from each other and floatdownstream on the top of the carrier material. After about 100 to 200seconds retention time, pellet identity is lost and a slag/clinker layeris formed on top of the molten carrier material. At this stage, it isvery important to prevent this slag/clinker layer lifting off thesurface of the molten carrier material thus precluding further reductiontaking place at an acceptably high rate.

Up to this point, no combustion is permitted above the floating chargepellet layer and, indeed, some cooling of roof refractories in this areaof the second hearth may be undertaken to delay pellet fusion.Thereafter, however, combustion of smelting reduction gases,principally, carbon monoxide to carbon dioxide, is encouraged to takeplace so that radiant energy is transmitted to the top of theslag/clinker surface and its temperature is progressively raised as thematerial floats away on top of the molten carrier material. Some 300seconds or so retention time in this fired region is typically requiredto bring the mean slag temperature up to around 1600° or 1650° C.,whilst the molten carrier material rises to no more than about 1500° C.having been introduced into this furnace hearth at about 1450° C.initially near where the ilmenite charge pellets are first admitted.This differential in temperature between molten carrier material and thefloating slag is in contradistinction to other smelting reductionprocesses using melt circulation, where the aim is to increase theemissivity of the surface of the molten carrier material without,however, introducing a large temperature drop across a very thin slaglayer. In the present case, the circulation rate and the furnacedimensions are preferably designed so that a titania slag layer,typically around 4 mm in thickness, is formed which, with postcombustion of CO to CO₂ in the freeboard above the melt, gives rise toradiative heat transfer intensities of around 125 kw/M², yieldingtemperature differences across the slag layer of about 160° C. Thus theobjective of generating high slag temperatures without having to exposethe whole melt circulation loop to excessively high temperatures isreadily obtained by this mechanism. This factor along with relativelylow melt velocities of about 5 cm/s in the first hearth enables a moltenlead hearth layer proposed for use in, for example, EP-A-0266975 to bedispensed with. Some localised cooling of the refractory in theimmediate area of the slag/liquid metal interface may also be introducedto prevent contamination of the slag product and to minimise refractoryattack.

At temperatures below about 1700° C. or so it is to be appreciated thatslags containing 90 percent TiO₂ or higher are not completely liquid butexist in a two phase region of liquid with a solid phase, which isreported in the literature to be rutile. Such two phase slags,particularly if some TiO₂ has been reduced to the lower oxide Ti₂ O₃,are known to possess very high viscosity, but nonetheless, they can beoverflown with the, molten carrier material, into a slag reservoir andaccumulated therein whilst the carrier material leaves this region viaan underflow weir. Whilst contained in this reservoir the slag may betop-blown with either a single oxygen lance or an array of lances topromote regeneration of TiO₂ from Ti₂ O₃ which releases sufficient heatto fuse the remaining solid rutile phase and provide enough surplus heatto satisfy the energy demands of some localised freeze-cooling at theslag/metal interface in this slag reservoir and to enable localisedcooling to generate a protective solid rutile layer on the walls of thereservoir itself. This exothermic heat generated in situ in the slag ismade available by purposely providing sufficient residence time in thesmelting reduction zone to convert a significant fraction of the TiO₂ toTi₂ O₃ and, depending on the exact chemical composition andconfiguration employed, this may be at least one half to two thirds ofthe TiO₂ in the feed concentrate being reduced initially to Ti₂ O₃. Thisreduction of TiO₂ is accomplished with dissolved carbon in the carriermelt and the dissolution requirements of fixed carbon from thecarbonaceous reductant in the second hearth of the melt circuit need tobe taken into account in providing adequate carbon dissolution for theprimary smelting reduction reactions and this additional servicerequirement.

During smelting reduction, a multitude of fine metal prills (droplets)are ejected into the slag layer and whilst the slag layer has a highviscosity it is difficult for these prills to settle back into the bulkiron-carbon melt. An important aspect of the operation of this titaniaslag reservoir is to ensure that high fluidity slag is generated and theopportunity is given for metal prills to settle out before the productslag is intermittently tapped or alternatively continuously overflowninto a receiving vessel or ladle. Typically, the slag leaves thereservoir at about 1730° C., whilst the molten carrier material leavesthe region normally around 1500° C., it being appreciated thatslag/carrier material contact area is minimised in this final part ofthe melt circuit so that such temperature differences are sustainable byvirtue of the relatively high circulation rates of metal in the circuitin terms of the actual metal production rate and the limited slag/metalinterfacial area provided when the slag reaches its highest temperaturelevel.

It will be appreciated that ilmenite smelting is more demanding in termsof carbon dissolution than say a normal smelting reduction process forcoal-based ironmaking. The necessity for having carbon levels not thatfar removed from carbon saturation as already discussed puts pressure onthe carbon dissolution area of the melt circulation loop. Despite this,it is possible with appropriate modifications, to retain the advantagesof using particulate coal (not pulverised) essentially in the asreceived condition with minimal preparation other than drying to removefree moisture with perhaps some coarse crushing and screening. To thisend, dry lump coal and associated fines can be added at one end of thesecond hearth so that the floating material lose their volatiles and thecoal is at least partly carbonised as it floats freely along withiron-carbon carrier material at a velocity in the region of say 20 cm/s.Under free floating conditions, a coal with an ash content of 8 percentcould be expected to generate (with flux addition to enhance fluidity ifnecessary) a coal ash slag layer typically less than 0.02 mm inthickness so that even quite fine coal particles floating with such alayer could be expected to penetrate into the iron-carbon meltthroughout most of their existence before liquid slag inhibits carbondissolution from their last remnants. The "hydraulic gradient"associated with the open-channel flow of the carrier ensures thatfloating coal ash slag does not accumulate other than at the far end ofthe second hearth where it is dammed to form a very shallow lake or poolof slag immediately upstream of an underflow weir at the downstream endof the second hearth. Typically this pool of slag is no more than a fewmillimeters or a centimeter in depth and its influence does not extendupstream significantly into the free flowing region. The purpose of thisdam is to constrain the lump carbonised coal so that it formseffectively an almost static layer of coked material in contact with theiron-carbon carrier material which flows virtually unimpeded underneaththe downstream weir to be removed continuously e.g, via an upleg of anRH vacuum lift and transported then into the first hearth. The principalmechanism of carbon dissolution under the conditions described is forcedconvection.

By controlling the rate of removal of the carrier material from underthe coked material (e.g, by controlling the flow of inert lift gas intothe RH upleg), it is possible to vary the melt flow rate according to apre-arranged schedule and in so doing, for example, if the flow isstopped momentarily or reduced appreciably, the slag pool upstream ofthe underflow weir and its associated residual coke particles and someiron-carbon melt will overflow the weir and be collected in a slagreservoir downstream of the weir. The coke particles in this reservoircan be gasified by a single top blow oxygen lance or, if moreconvenient, an array of such oxygen lances, such that eventually slagoverflowing as discharged from this reservoir contains very littleresidual unburnt coke. The CO/CO₂ gases generated pass back up thesecond hearth to be combusted above the melt along with coal volatilesso that the heat thus generated satisfies not only the coalcarbonisation and dissolution requirements, but also is picked up by thecarrier iron-carbon melt so that heat is transferred as sensible heat inthe melt to satisfy the thermal demands of smelting reduction in thefirst hearth. The associated iron-carbon melt that also overflowsintermittently separates out and then rejoins the principal metal flowvia which has passed through the underflow weir.

The total coal requirement for smelting one tonne of high-grade ilmeniteconcentrate containing 63% TiO₂ and 32% ferric oxide to a titaniferousslag with a composition typically of 90% TiO₂ and about 5% ferrous oxidealong with molten iron containing about 4% carbon is calculated to bearound 0.2 tonne of medium volatile coal with the following analysis(dry basis):

    ______________________________________                                                                              %                                                                        %    Fixed  %                                % C   % H    % N      %5   % O   Ash  C      Volatile                         ______________________________________                                        63.7  4.83   1.4      0.99 1.8   7.22 71.7   21.0                             ______________________________________                                         Calculated Gross Calorific Value = 33,346 kJ/kg                          

It needs a projected area of about 185 m² of coked coal melt interfaceto attain the fixed carbon dissolution rate required to produce 100,000tonnes per year of titanium dioxide contained in a slag of 90% TiO₂content. Assuming that the second hearth is 7.5 m wide and 35 m inoverall length, some 25 m of this length is committed to the maintenanceof the virtually static floating layer of coked coal held in place bythe downstream underflow weir. With iron-carbon melt flowing underneaththis coked coal layer at a velocity of about 20 cm/s, the required fixedcarbon dissolution rate can be sustained with an iron-carbon meltcirculation rate of about 45 tonne per minute which is about thecirculation rate attained by a moderately sized RH steel degassing unitin a commercial practice. On the same basis the size of the first hearthwould be about 5 m wide with an overall length of about 30 m.

For the present example it can be shown that the thermal energyrequirement per tonne of products (titania slag plus hot metal product)is about 10 GJ/per tonne, whereas it is believed that somewhere in theregion of 25 GJ per tonne of products is required for the existingelectric furnace technology as practised, for example, at Richards Bayin South Africa, assuming of course that the electricity being used isthermally generated from fossil fuel.

Preferably, the FeO content of the titania enriched slag is not morethan 5% by weight and the titania content is not less than about 90% byweight.

Accurate temperature control is also important in preventing excessiveover-reduction of TiO₂. Temperature control in the process of thepresent invention is simplified because it is possible to maintain avirtually uniform temperature, or at least minimal temperaturevariation, throughout the whole of the circuit because of the "heatsink" property of the large mass of circulating molten carrier material.

Carbon monoxide, hydrogen and other coal volatiles produced as a resultof carbonisation of the carbonaceous reductant can be burnt in the firsthearth whilst carbon monoxide produced as a result of the reduction ofiron oxide in the titaniferous material and the partial reduction ofTiO₂ to Ti₂ O₃ can be burnt in the second hearth, to form carbon dioxidein both hearths since it is not normally practicable to transportextremely hot gases between the hearths. In a preferred arrangement,accumulation of slag in both hearths is preventeded by continuouslyremoving the slag from the hearths in such a manner that the heat ofsuch combustion is transferred effectively to the relatively clean meltsurface in the second hearth, but in the first hearth a somewhat thickerlayer of slag (typically about 4 mm thick) is formed whereby the titaniaslag temperature is raised appreciably above that of the underlyingmetal.

BRIEF DESCRIPTION OF THE DRAWINGS

An embodiment of the present invention will be now be described, by wayof example, with reference to the accompanying drawing which is aschematic plan view of equipment for the production of hot metal, ie, Fe(C) and high purity TiO₂ slag.

Referring now to the drawing, the plant comprises a first or upperhearth 10, a second or lower hearth 11, a high purity TiO₂ slag removalchamber or slag reservoir 12, a carbonaceous reductant feed chamber 14,a coal ash slag removal chamber 16, and a chamber 18.

The upper hearth 10 discharges via an overflow weir 20 into the chamber12 which is connected with chamber 14 via an underflow weir 22. Thechamber 14 discharges into an upstream end of the lower hearth 11 via anoverflow weir 24. The downstream end of the lower hearth 11 is connectedwith chamber 16 via an underflow weir 26. The chamber 16 communicateswith the chamber 18 via an R-H unit 28 having an inlet snorkel 30 in thechamber 16 and an outlet snorkel 32 in the chamber 18.

In use, operation of the R-H unit 28 serves to circulate molten carriermaterial (iron containing dissolved carbon) at a speed of about 5 cm/sin a closed loop extraction circuit serially through chamber 18, upperhearth 10, chamber 12, chamber 14, lower hearth 11 and chamber 16. Drylump coal (and associated fines) as the carbonaceous reductant isintroduced into the chamber 14 via line 34 on top of the molten carriermaterial before flowing over weir 24 into the upstream end of the lowerhearth 11. During passage of the molten carrier material and coal alongthe lower hearth 11, the coal carbonises to form carbon, carbonmonoxide, hydrogen and other coal volatiles. Coal ash slag forms on topof the molten carrier material in the lower hearth 11 as a moving thinlayer (typically less than 0.02 mm thick). Such coal ash slag is dammedby the underflow weir 26 at the downstream end of the lower hearth 11 sothat it collects as a stationary layer at the downstream end region ofthe hearth 11. This causes a relatively stationary layer of coked coalto accumulate at the downstream end region of the hearth 11. Meanwhile,the molten carrier material continues to pass under the relativelystationary layer of coked coal and thereby picks up carbon therefrom.The addition of the coal and the temperature are closely controlled soas to ensure that the carbon content of the hot metal forming the moltencarrier material is in the range of 3 to 5% by weight.

The molten carrier material containing the dissolved carbon is removedfrom under the slag in the chamber 16 via the inlet snorkel 30 of theR-H unit 28. Periodically, the removal rate via snorkel 30 is slowed tocause accumulated coal ash slag and associated coke particles and someof the carrier material to overflow the weir 26 and collect in thechamber 16 where the coke particles are gasified by one or more top blowoxygen lances 35. The resultant coal ash slag is removed from chamber 16via line 36. The CO/CO₂ gases generated by the top blowing pass back upthe second hearth 11 to be combusted with air via lines 47 above themelt along with the coal volatiles so that the heat thus generatedsatisfies not only the coal carbonisation and dissolution requirement,but also is picked up by the relatively clean carrier material towardsthe upstream end of the hearth 11 so that heat is transferred assensible heat to assist in satisfying the thermal demands of thesmelting reduction in the first hearth 10.

The molten carrier material which is now free of the coal ash slag istransported into chamber 18 via the R-H unit 28 and passes from there tothe hearth 10 via channel 40. Cold pellets of titaniferous material, egilmenite concentrate or titaniferous magnetite, are introduced via feed38 extending across the hearth 10 gently onto the carrier material afterthe latter has passed through the channel 40. As described hereinbefore,the pellets of titaniferous material are caused to float on the surfaceof the carrier material for an extended period whilst being essentiallyunheated from above, during which time the iron oxide in thetitaniferous material reacts with the carbon in the carrier material toform iron and carbon monoxide and some of the TiO₂ is reduced to Ti₂ O₃.The material is then heated from above by combusting the carbon monoxidewith preheated air via lines 46 to produce a partially fused layer oftitania slag containing Ti₂ O₃ (about 4 mm thick) on top of the carriermaterial in hearth 10. This slag 20 together with the molten carriermaterial overflows via weir 20 into the chamber 12 where the slag is topblown with oxygen via line 41 to convert Ti₂ O₃ to TiO₂ and to releaseheat which ensures that the slag is in a sufficiently fluid state torelease entrained metal prills. Localised cooling is effected in chamber12 as described previously herein. The resultant high purity titaniaslag product (typically about 90% titania and 5% FeO) is removed vialine 42. The molten carrier material from which the high purity TiO₂slag has been removed then flows via underflow weir 22 into the chamber14.

Hot metal is bled off from the circuit via line 44 which is located inthe chamber 14 immediately adjacent the underflow weir 22, although itmay be removed at any convenient location in the circuit.

Instead of being separate therefrom, the chambers 12 and 18 may beformed by regions of the first hearth 10 divided from the remainder ofthe latter by respective weirs, and likewise for the chambers 14 and 16in respect of the hearth 11. Thus, the chambers 12 and 18, even thoughthey are shown as separate items, can be considered to constitutedownstream and upstream end regions, respectively, of the first hearth10. Likewise, the chambers 14 and 16 can be considered to constituteupstream and downstream end regions, respectively, of the second hearth11.

I claim:
 1. A method of upgrading titaniferous material; containing ironoxide, comprising the steps of circulating a molten carrier materialcomprising molten iron and dissolved carbon in a closed loop paththrough first and second hearths; introducing the titaniferous materialinto the carrier material; heating the introduced titaniferous materialin a heating zone in the first hearth so as to reduce iron oxide to ironand produce a titania slag having a solid rutile phase; removing thetitania enriched slag from the molten carrier material before the latteris circulated to the second hearth; introducing carbonaceous reductantinto the molten carrier material so as to cause carbon to dissolve inthe carrier in the second hearth; and performing a further slag removaloperation on the circulating molten carrier material before passing thelatter to the first hearth.
 2. A method as claimed in claim 1, whereinthe amount of dissolved carbon in the molten iron contacted by thetitaniferous material exceeds 2% by weight.
 3. A method as claimed inclaim 1, wherein the amount of dissolved carbon is 3 to 5% by weight. 4.A method as claimed in claim 1, wherein the titania enriched slagcontains not more than 5% by weight of Fe and the titanium content isnot less than about 90% by weight.
 5. A method as claimed in claim 1,wherein the titaniferous material is introduced in the form of pelletsonto the surface of the molten iron containing dissolved carbon.
 6. Amethod as claimed in claim 1, wherein the molten iron containingdissolved carbon is maintained at a lower temperature than the titaniaenriched slag in contact therewith.
 7. A method as claimed in claim 1,wherein the titaniferous material is introduced so as to form a layer onthe molten iron containing dissolved carbon and is heated from aboveafter a delay upon introduction so as to delay fusion of thetitaniferous material.
 8. A method as claimed in claim 1 wherein thetitaniferous material contains TiO₂, a proportion of which is reduced toTi₂ O₃ during reduction of the iron oxide to iron so that the titaniaslag contains a proportion of Ti₂ O₃, such that slag is passed to a slagreservoir, and the Ti₂ O₃ is oxidized to TiO₂ so that sufficient heat isreleased in the slag reservoir to fuse any remaining solid rutile phaseIn the slag before the slag product is removed.
 9. A method as claimedin claim 1, wherein the carbonaceous reductant is introduced into thecarrier material by establishing a relatively stationary layer of thecarbonaceous reductant on the carrier material and causing the carriermaterial to flow under such layer whereby to pick up carbon from suchlayer of reductant.